Production of copper via looping oxidation process

ABSTRACT

Copper is produced by a looping oxidizing process wherein oxidation of copper sulfide concentrate to molten blister copper by conversion with copper oxides (and optionally oxygen from air) in a one step, molten bath operation to produce molten blister copper, iron oxide slag, and rich SO 2  off gas. The blister copper is treated in an anode furnace to reduce the iron content and oxidize residual sulfur, and prepare it for either electrolysis or reoxidation.

CROSS-REFERENCE TO RELATED APPLICATIONS AND PUBLICATIONS

This application claims priority from U.S. provisional patentapplication 61/662,603 and 61/690,210 both filed Jun. 21, 2012. The fullcontent of the said applications and of all other patents, publishedpatent applications and non-patent publications cited herein areincorporated herein by reference as though set out at length herein.

FIELD OF THE INVENTION

The present invention relates to improved methods for production ofcopper from copper sulfide concentrates produced as part of a mineralore refining.

BACKGROUND OF THE INVENTION

De Re Metallica by Georgius Agricola, published in 1556, details themining, smelting, and refining techniques and technologies of that era.Since then the basic chemical reactions to produce copper have notsignificantly changed, while the modern smelting process now treats aconcentrate rather than as-mined ore of that time. However, technologyhas markedly advanced through numerous changes and improvements tocopper smelting methodology since De Re Metallica's publication. The“Welsh” process, based on a series of sequential reverberatory smeltingsteps, subsequently dominated copper smelting for over a hundred years.In the 1890s, Nicholls and James developed a process (Great BritainPatent 18,898) based on an alternative final step in the traditional“Welsh” copper smelting process. In this invention part of thehigh-grade white metal stream was diverted for calcination to produce acopper oxide material for subsequent re-use in the oxidation of the mainwhite metal stream to produce metallic copper. The large, fuel-firedreverberatory furnace was later used for concentrate smelting throughoutthe first three-quarters of the twentieth century. In more modern times,newer flash and bath smelting processes were developed. The flashsmelting concept was described by Bryk et al. in U.S. Pat. No.2,506,557. Later, Gordon et al described a variant of the flash smeltingprocess in U.S. Pat. No. 2,668,107. An alternative to flash smelting isthe bath smelting process such as introduced by McKerrow et al. in U.S.Pat. No. 4,005,856 and also Bailey et al. in U.S. Pat. No. 4,504,309.Still another bath smelting approach, referred to as the Isasmeltprocess, based on a top lance blowing system with the particular lancesystem described by Floyd in U.S. Pat. Nos. 3,905,807 and 4,251,271, wasdeveloped. The lance system is used in the process operating in Arizonaas described by Bhappu et al in: EPD Congress 1994, Edited by G. Warren,The Minerals, Metals and Materials Society, 1993, pages 555 to 570. Eachof the contemporary processes described above for the modern era producea medium to high-grade of copper matte which is typically processed inPeirce-Smith converters to blister copper. Following this, the producedcopper is transferred to an anode furnace (European Patent 0648849 B2)for finishing to anode copper for subsequent casting and thence toelectrolytic refining. The conventional flash furnace and converterprocess flow sheet is depicted in FIG. 1. As shown here, copperconcentrate is introduced into the flash smelting furnace (as an exampleof a modern smelting unit) where the copper sulfide concentrate reactwith oxygen-enriched air to form a medium grade of matte and a slag. Thereaction in the flash furnace can be represented by the followingequation (Equation 1). Some nitrogen will also be present with theoxygen, depending on the degree of oxygen enrichment.

$\begin{matrix}{{{2\; {CuFeS}_{2}} + {\frac{13}{4}O_{2}}} = {{{{Cu}_{2}S} + {\frac{1}{2}{FeS}} + {\frac{3}{2}{FeO}} + {\frac{5}{2}{SO}_{2}\Delta \; H^{o}}} = {{- 250}\mspace{14mu} {Wh}}}} & (1)\end{matrix}$

A fossil fuel may be used as a supplementary energy source as requiredfor heating/sustaining typical flash temperatures above 1350° C. Asilica flux is added during this step to flux with the iron oxideproduct shown in Equation (1). The resulting flash furnace slag is sentto a slag treatment facility for copper recovery. The process off-gasesare first cleaned and are then treated in a sulfuric acid plant forsulfur recovery.

The remaining molten white metal is transferred to a converter, where itis blasted with oxygen-enriched air to remove remaining sulfides,produce the blister copper, and form an additional slag (Equations 2 and3).

Cu₂S+O₂=2Cu+SO₂ ΔH°=−59 Wh  (2)

FeS+1.5O₂=FeO+SO₂ ΔH°=−130 Wh  (3)

The converter slag is typically higher in copper content, and alsorequires slag treatment. The flue gases from this step also requireprocessing in the sulfuric acid plant. The copper melt is sent to anodecasting (often proceeded by an anode furnace to further purify thecopper metal) and then on to electrolysis.

In total, this flash process has gained wide-spread acceptance in thecopper industry. Its advantages over older reverberatory molten bathsmelting are manifold: utilization of the heat released during oxidationof sulfides with oxygen, high furnace throughput, high copper recoveryinto matte, and higher SO₂ content in the off gas relative to the moltenbath process. However, and as previously mentioned, significant controlmust be maintained throughout the process and significant opportunitiesfor improvement exist. Principally, the composition of the feedmaterials must be well specified, an understanding of the absolute andrelative particle sizes is required, moisture and sulfide contents ofthe concentrates and fluxes must be quantitatively known, and furnacedimensions and temperatures are critical. Precise control over the feedratios and rate of oxygen injection must be maintained. Similarly, theamount of siliceous flux that must be added is wholly dependent on thesulfide concentrate and the amount of iron that must be oxidized; highcopper losses into the slag are still observed and this requires aseparate treatment step. The energy demands of the flash process requirepreheating of the furnace to circa 900-1100° C. to initiate theexothermic reactions involved when oxygen enrichment is not used. Thishigh temperature conversion leads to NO, formation. Oxygen-enriched airis normally used, in which case preheating the air is not common.

Several variations on flash smelting technology have been developedsince the Gordon et al. first work. U.S. Pat. Nos. 5,662,730; 3,790,366;3,948,639; 3,892,560; 4,615,729; 4,470,845; 3,674,463; 5,607,495;4,521,245; and US Published Patent Application 2005/0199095 demonstrateoxygen enrichment of air, various techniques for copper recovery fromslags as well as partial or dead roasting of the sulfide concentrateprior to flash smelting.

Work performed in the 1890s by Thomas Davies Nicholls, et al. (GreatBritain patent 18,898) details the use of copper oxides in roastingcopper mattes to copper metal. During this time period, pneumatic copperconverting was just in its infancy, hence this method was considered animprovement over the established contemporary roasting process. Copper(I) sulfide, previously smelted into matte (76-78% copper), is crushedand melted in a reverberatory furnace common at that time with calcinedcopper. The produced copper was then poled to produce a final copper. Inthis process, it was difficult to produce CuO during the calcination ofCu metal, so Cu₂O was used. Production of copper anodes from coppersulfide sources without producing an intermediate copper matte phase hasbeen performed and summarized in the literature^(1,2). In suchoperations, the copper sulfide concentrate is first dead roasted atelevated temperatures (900° C.) in an excess of oxygen to produce acopper calcine with sulfur levels around 2% (generally 1-1.5% sulfur).The calcine is then transferred to an electric furnace (e.g. theBrixlegg Process)^(3,4), a segregation furnaces^(5,6), a rotaryfurnace⁷, or a shaft furnace^(8,9) where it is further converted toproduce blister copper, slag and SO₂ off gases. ¹Opie W R, (1981)Pyrometallurgical processes that produce blister grade copper withoutmatte smelting. IMM, 137-140.² (1980) Dead Roast-Shaft Furnace coppersmelting, World Mining, Vol 33, Issue 12, 40-41.³ Kettner P, Maelzer CA, and Schwartz W H, (1972) The Brixlegg Electro-Smelting ProcessApplied to Copper Concentrates, AIME Annual Meeting, San Francisco.⁴Paulson D L, Worthington R B, and Hunter W L, (1976) Production ofBlister Copper by Electric Furnace Smelting of Dead-Burned CopperSulfide Concentrates, U.S. Bureau of Mines, RI-8131.⁵ Opie W R, andCoffin L D, (1974) Roasting of Copper Sulfide Concentrates Combined withSolid State Segregation Reduction to Recover Copper, U.S. Pat. No.3,799,764.⁶ Pinkney E T, and Plint N, (1968) Treatment of RefractoryCopper Ores by the Segregation Process, Transactions of AIME, Vol 241,373-415.⁷Rajcevic H P, Opie W R, and Cusanelli D C (1978) Production ofBlister Copper in a Rotary Furnace from Calcined Copper-IronConcentrates, U.S. Pat. No. 4,072,507.⁸Rajcevic H P, Opie W R, andCusanelli D C (1977) Production of Blister Copper Directly from DeadRoasted-Copper-Iron Concentrates Using a Shallow Bed Reactor, U.S. Pat.No. 4,006,010.⁹ Opie W R, Rajcevic H P, Querijero E R, (1979) DeadRoasting and Blast-Furnace Smelting of Chalcopyrite Concentrates,Journal of Metals, Vol 31, Issue 7, 17-22.

It is an object of the present invention to provide a better method torecover copper from copper sulfide concentrates via a process chemistrypreviously unused by the copper smelting industry. This process isreferred to as the “Looping Sulfide Oxidation” (or “LSO”) process.

SUMMARY OF THE INVENTION

Many of the opportunities for improvement in flash smelting outlinedabove stem from the incremental removal of sulfur in two separateprocessing steps. As a result, the concentrations of the SO₂ streams,which while higher than the concentrations in the roaster andreverberatory furnace off gases (ca. 15-20% SO₂), the presence of twosulfurous off gas streams requires handling and treatment, and slagswith relatively high copper contents are produced in both the flashfurnace and the converter. The Looping Sulfide Oxidation process forcopper production removes sulfur in a single step while using copperoxides (Cu₂O and CuO) as oxidizing agents to either replace or augmentoxygen (O₂) from natural air without producing a matte phase. Referenceherein to copper oxide oxidizing agents include copper carbonates,sulfates and other oxygen containing copper compounds thermodynamicallysuitable for use in the Looping Sulfide Oxidation process following theguidelines shown in this application.

Looping Sulfide Oxidation features three distinct steps: conversion ofthe copper sulfide concentrates into copper and copper oxides (wholesaledesulfurization), recovery of copper from the slag, and looping oxideregeneration (FIG. 2). This process primarily uses CuO as the oxidizingagent instead of O₂ in order to eliminate oxygen-enriched airutilization in the sulfur removal step and to generate energy from thereoxidation of copper downstream. Looping Sulfide Oxidation allows forgreater energy capture by performing all the desulfurization ofconcentrates in a single step. Metal refining and slag treatment arehandled simultaneously in the second step. Overall copper yield matcheswell with recovery levels achieved in the conventional flash process.

In this first step of the Looping Sulfide Oxidation process, the copperconcentrate is blended with fluxes and the oxidizing agent, CuO. Inalternative embodiments, the CuO may be augmented with oxygen from airin a fashion such that the total stoichiometry of the system ismaintained. The reaction that takes place in this furnace is presentedbelow.

$\begin{matrix}\left. {{CuFeS}_{2} + {aCuO} + {\left( \frac{5 - a}{2} \right)O_{2}}}\rightarrow{{\left( {1 + a} \right){Cu}} + {FeO} + {2\; {SO}_{2}}} \right. & (4)\end{matrix}$

In such a reaction scheme, the value of a is allowed to vary such thatratio of CuO to O₂ might range from 5:0 to minimal CuO with greaterportions of O₂ while still satisfying the reaction stoichiometry. Whilethe relative ratio of CuO and O₂ is important, the total amount ofoxidizer may be equal to or in excess of the amount required tocompletely oxidize the copper concentrate. Consideration must be madethat excess of the oxidizer can influence the copper melt and/or slagcompositions. In this sense, CuO functions to oxidize the iron in theconcentrate and/or slag in addition to oxidizing (desulfurizing) thecopper in the concentrate. A fraction of copper will be present in theslag as Cu₂O due to the equilibrium established between the slag and thecopper metal phase. As such, the calculated stoichiometry of theoxidizing agents is minimal, and will be exceeded.

One possible embodiment of the furnace is a Vanyukov-typefurnace^(10,11) (exemplars of which appear in U.S. Pat. Nos. 4,252,560and 4,294,433), i.e. the concentrate and fluxes are added through theslag, which is agitated by the injection of N₂, hot combustion products,and/or air through tuyères; additionally, the energy is supplied viaelectrodes submerged in the slag. Due to the high energy demand of theendothermic reaction that takes place, additional heat must be providedto the first furnace. This heat will be supplied either solely throughthe electrical heating of the furnace or through electrical heatingaugmented by combustion of fuels, whose heat will be transmitted to thefurnace through the hot gases in the tuyères and whose chemically inertcombustion product gases will be injected into the molten slag tofacilitate mixing. Another embodiment may use a top-blown lance in theslag in an Isasmelt-type furnace; this embodiment may also includeelectrode heating. ¹⁰Bystov, V P, Fyodorov, A N, Komkov A A, and SorokinM L (1992) The use of the Vanyukov process for the smelting of variouscharges, in Extractive Metallurgy of Gold and Base Metals, AustralasianInstitute of Mines and Metallurgy, Pardville, Vic., 477-482.¹¹Bystrov, VP, Komkov A A, and Smimov L A (1995) Optimizing the Vanyukov process andfurnace for treatment of complex copper charges, in Copper 95-Cobre 95,Vol. IV—Pyrometallurgy of Copper, ed. Chen W J, Diaz C, Luraschi A, andMackey P J, The Metallurgical Society of CIM, Montreal, Canada, 167-178.

The process chemistry that takes place in the first furnace is ofcritical importance. Most notably, metal not matte is formed during thisstep. This marks a significant differentiation and improvement over thepresent state of the art. The molten copper metal produced in thefurnace is very low in iron and is sent directly to the anode furnace.The oxidized iron slag contains copper that must be recovered duringslag treatment. As previously mentioned, the complete desulfurization ofthe concentrate is accomplished in this single step. This allows forsignificant energy capture during sulfuric acid production in an acidplant. Additionally, because no sulfurous/sulfuric gases will beproduced in the downstream processing, aggressive energy capture can beperformed on the off gases without fear of acid condensation. The majordifferences between this invention and the closest prior art (Nichollset al.) are:

-   -   1. The raw material is neither blister copper nor matte, but is        rather copper sulfide concentrate    -   2. The raw material and copper oxide are simultaneously fed into        the molten slag in the smelting furnace; the slag is agitated        via the injection of combustion product gases or chemically        inert gases    -   3. The use of oxygen is controlled and special care is taken to        ensure that the total oxygen from air and copper oxide does not        exceed 20% excess of the required stoichiometric amounts.        Slag composition in the smelting step can be further optimized        by changing the amounts of fluxes (CaO, SiO₂, Al₂O₃) added to        reduce the viscosity, lower the melting temperature, and        increase copper recovery into the copper melt. Increased calcium        oxide will decrease the copper solubility in the slag. The        amount of Fe₂O₃ (i.e. the amount of Fe³⁺) in the slag must be        reduced.

During slag treatment, the goal is to recover as much copper as possiblefrom the slag phase so that it can be returned to the processing loopfor copper anode production. In general, the slag from the first furnacewill contain ca. 10-15% copper in the slag as Cu₂O. The slag, which isstill molten, is treated with either carbon (from coal or natural gas)to reduce the copper oxides to copper metal (and the trivalent iron todivalent iron), or oxidized with sulfur (e.g., as iron pyrite), toproduce copper matte. With carbon reduction the copper from the slagtreatment furnace can be mixed with the copper rich material from thesmelting furnace; with sulfidation, the matte will be returned to thesmelting furnace to be reprocessed.

Slag treatment must reduce the copper content in the waste slag tolevels below ca. 0.4 weight percent. The copper solubility in the slagis a function of many variables; one of critical importance is theFe(III):Fe(II) ratio. In this process, the copper solubility in the slagis reduced (and thereby the copper recovery is increased) bysignificantly reducing the Fe(III) content in the slag. Additionally,when the product from slag treatment is copper metal, the iron contentmust be sufficiently low enough for an anode furnace. In the processpresented here, the copper metal from the slag treatment step is blendedwith the copper metal from the first furnace to produce a copper-richstream to be processed in the anode furnace. If sulfidation isperformed, the copper matte produced will be processed in the firstfurnace. This step in the process is carried out in a traditional slagtreatment furnace, e.g. an electric furnace.

The anode furnace operates in the same fashion as conventional anodefurnaces. The copper melt is first oxidized to oxidize any residual ironto a dry slag; in this step some of the copper metal may be co-oxidized.The slag is tapped off and the remaining copper melt is then deoxidizedprior to casting to anodes ready for electrolytic refining.

The fraction of copper that is sent to electrolysis is determined by thestoichiometry of the reaction in the smelting furnace (i.e. the amountof copper in the concentrate is equal to the amount of copper in theanodes for electrolysis). The necessary amount of copper to produce therequisite copper oxide for oxidation of the copper concentrate is sentto the reoxidation furnace. In this furnace, the copper melt is atomizedand oxidized to CuO with air. This highly exothermic reaction can beharnessed for energy capture. The molten copper is oxidized at hightemperatures in a downer or vertical furnace (ca. 1500° C.), and cooledbelow freezing to ca. 800° C. The powdered CuO is then looped back tothe smelting furnace to complete the reaction cycle.

2Cu+O₂→2CuO ΔH_(1500° C.)=−59.5 Wh  (5)

This invention provides an improvement over the closest prior artwherein Cu₂O was produced (Nicholls et al.) in which copper matte isoxidized to produce copper oxide. In this work, copper is reoxidizedafter atomization to promote rapid and complete oxidation.

Alternatively, other sources of copper oxides can be used as oxygencarriers during Looping Sulfide Oxidation. For example, CuO is used inthe industry as pigments in ceramic materials, battery materials andcatalysts. These materials can be fed to the smelting furnace to augmentthe copper oxides that are produced in the reoxidation furnace.Similarly, several copper oxide minerals are processed by the copperindustry; these minerals can be used as source of copper oxides duringLooping Sulfide Oxidation. Thermodynamic calculations, made withFactSage 6.4¹² thermodynamic software, detailing such operation aredisclosed below. ¹²Bale, C. W., et al., FactSage™ 6.4.1, Thermfact andGTT-Technologies, CRCT, Montreal, Canada (2013).

Copper scrap is also an important copper stream for Looping SulfideOxidation. Copper scrap metals and copper alloy scrap can be processedin Looping Sulfide Oxidation via either smelting in the smelting furnacein the presence of copper oxides (potentially augmented with air), orvia initial oxidation to copper oxides in the reoxidation furnace. Inthe former embodiment, the copper scrap is melted in the smeltingfurnace and converted to copper metal in the same fashion as copperconcentrate. Depending on the composition of the scrap, alloyed metalswill report to either the slag or the copper phase. The use of thisembodiment can gain an increase in the iron content in the molten copperdue to the reduction of the iron oxides present in the slag with anyreducing metals (e.g. aluminum or silicon) present in the scrap. In thelatter embodiment, the copper scrap is processed to enable its rapidatomization and oxidation (in one embodiment, in a plasma furnace) tocopper oxides that can be looped to the smelting furnace.

Other objects, features and advantages will be apparent from thefollowing detailed description of preferred embodiments taken inconjunction with the accompanying drawings in which:

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 (Prior Art) shows in block diagram form a generalized processflow chart for flash smelting conversion;

FIG. 2 shows in block diagram form a Looping Sulfide Oxidation Processto produce anode copper;

FIG. 2 a shows schematically an electric arc furnace used in thesmelting conversion;

FIGS. 3-8 are traces of thermodynamic data showing calculations ofproduction conditions (CuO) feed variation on output conditions of thecopper melt and slag during the smelting step;

FIGS. 9-15 show traces of thermodynamic data detailing the slagtreatment and output of the slag treatment furnace, the treated slag andthe copper melt or copper matte;

FIGS. 16-19 show traces of thermodynamic data detailing the smelting ofCuFeS₂ with CuCO₃; and

FIGS. 20-22 show traces of thermodynamic data detailing the smelting ofCuFeS₂ with CuSO₄.

DETAILED DESCRIPTION OF PREFERRED EMBODIMENTS Example 1

In the present analytical example (not based on a physical plantactually constructed) the process is described on a production basis ofapproximately 1000 kg of anode copper. The process flow (all or parts ofwhich can be continuous, semi-continuous or batch format) is shown inFIG. 2 and the preferred basic configuration of the electric furnace (anarc furnace) is shown in FIG. 2 a including tuyères for gas injectioninto a molten slag formed in the furnace.

Electric Furnace

A room temperature copper concentrate comprising 3000 kg CuFeS₂, 173.4kg FeS₂, and 294.8 kg gangue (CaO, Al₂O₃, SiO₂), preferably in freeflowing powder form, is to be mixed with 7400 kg of CuO at 800° C. inthe first smelting furnace (Table 1). Heat and material balances werecalculated using HSC 7.1 Chemistry for Windows thermochemicalsoftware¹³. Silica (1000 kg) and lime (500 kg) fluxes are also taken asto be added to the melt. The melt is to be heated to 1300° C. viaelectrical and/or combustion heating. The reaction produces a metalliccopper melt, an oxidized slag, and a rich SO₂ gas stream. In thisExample, 14% excess CuO is used to produce an optimal copper melt and anoptimal slag (FIGS. 3, 4, and 5). The copper melt is 98.8% copper with0.002% Fe, and 0.88% S (FIG. 6). The slag includes some copper oxide (asCu₂O), iron oxides and gangue and flux derivatives. All compositionsherein are weight percent unless otherwise noted. ¹³ Roine, A., et al.,HSC 7.11, Outotec, Pori, Finland (2011).

As discussed in the above Summary of the Invention, the coppersolubility in the slag is largely dependent on the degree of oxidationof the iron also present in the slag. The fluxes added to the furnaceare designed to aid in slag formation and produce a low melting, fluidslag. The slag produced in this Example melts at 110° C. with aviscosity of 2.0 poise (at 1300° C.). The Cu₂O content in the slag is13.2%, and requires treatment to recover as much of this copper aspossible (FIG. 7). FIG. 7 demonstrates that during smelting, the coppercontent in the slag is largely independent of the slag composition andoperating temperature. However, as shown in comparing FIGS. 8 and 9, thedramatically higher O₂ partial pressure above the slag in the electricfurnace as compared to the O₂ partial pressure above the slag in theslag treatment furnace leads to different slag chemistries. Mostnotably, the decreased copper solubility in the slag after slagtreatment can be explained by considering the lower oxygen partialpressure present in the treatment furnace. This demonstrates that thecopper content in the slag can be controlled by the oxygen partialpressure.

TABLE 1 Electric Furnace Heat & Material Balance Temperature Amount,Amount, Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUT CuConcentrate 25.000 18.853 3468.198 0.855 0.00 −2180.49 CuFeS2 25.00016.348 3000.000 0.714 0.00 −864.48 FeS2 25.000 1.445 173.400 0.035 0.00−71.56 CaO*Al203*2SiO2 25.000 1.060 294.798 0.107 0.00 −1244.44 RecycledCuO 800.000 93.029 7400.000 1.173 1025.04 −3040.27 CuO 800.000 93.0297400.000 1.173 1025.04 −3040.27 Flux 25.000 25.559 1500.000 0.534 0.00−5783.49 SiO2 25.000 16.643 1000.000 0.385 0.00 −4211.01 CaO 25.0008.916 500.000 0.150 0.00 −1572.47 Heating 25.000 47.972 1243.806 888.6580.00 0.00 C 25.000 8.326 100.000 0.044 0.00 0.00 O2(g) 25.000 8.326266.412 186.609 0.00 0.00 N2(g) 25.000 31.320 877.394 702.005 0.00 0.00Energy Required 2493.56 OUTPUT Copper Melt 1300.000 105.915 6607.98731.555 1449.45 1544.56 Cu 1300.000 102.724 6527.713 0.729 1417.281417.28 Fe 1300.000 0.003 0.142 0.000 0.03 0.03 S 1300.000 1.814 58.1470.028 21.91 21.91 O(g) 1300.000 1.374 21.985 30.799 10.22 105.33 Slag1300.000 47.844 3596.867 0.940 1233.75 −7504.18 Al2O3 1300.000 1.059108.020 0.027 44.85 −448.28 SiO2 1300.000 18.759 1127.114 0.434 461.01−4285.28 CaO 1300.000 9.973 559.294 0.167 178.67 −1580.27 FeO 1300.00011.671 838.491 0.140 240.41 −641.78 Fe2O3 1300.000 3.057 488.235 0.093154.03 −547.63 Cu2O 1300.000 3.325 475.714 0.079 154.79 −0.93 Flue Gas1300.000 73.419 3407.221 1645.574 1143.46 −2551.06 SO2(g) 1300.00033.772 2163.415 756.960 634.42 −2150.05 CO2(g) 1300.000 8.326 366.412186.609 152.67 −757.38 N2(g) 1300.000 31.320 877.394 702.005 356.37356.37

TABLE 2 Slag Treatment Furnace Heat & Material Balance TemperatureAmount, Amount, Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUTSlag Furnace Slag 1300.000 47.844 3596.867 0.940 1233.75 −7504.18 Al2O31300.000 1.059 108.020 0.027 44.85 −448.28 SiO2 1300.000 18.759 1127.1140.434 461.01 −4285.28 CaO 1300.000 9.973 559.294 0.167 178.67 −1580.27FeO 1300.000 11.671 838.491 0.140 240.41 −641.78 Fe2O3 1300.000 3.057488.235 0.093 154.03 −547.63 Cu2O 1300.000 3.325 475.714 0.079 154.79-0.93 Reductant 25.000 4.329 52.000 0.023 0.00 0.00 C 25.000 4.32952.000 0.023 0.00 0.00 S 25.000 0.000 0.000 0.000 0.00 0.00 EnergyRequired 76.99 OUTPUT Copper Melt 1300.000 6.573 417.129 0.053 90.6790.68 Cu 1300.000 6.498 412.946 0.046 89.66 89.66 Fe 1300.000 0.0754.179 0.001 1.01 1.01 O(g) 1300.000 0.000 0.004 0.006 0.00 0.02 TreatedSlag 1300.000 47.430 3080.021 0.843 1054.22 −7298.04 Al2O3 1300.0001.059 108.019 0.027 44.85 −448.28 SiO2 1300.000 18.759 1127.126 0.434461.01 −4285.33 CaO 1300.000 9.973 559.297 0.167 178.68 −1580.28 FeO1300.000 17.416 1251.250 0.209 358.75 −957.71 Fe2O3 1300.000 0.14723.551 0.004 7.43 −26.42 Cu2O 1300.000 0.075 10.778 0.002 3.51 −0.02Flue Gas 1300.000 4.330 151.739 97.040 62.79 −219.84 CO(g) 1300.0002.425 67.932 54.358 27.86 −46.60 CO2(g) 1300.000 1.904 83.807 42.68234.92 −173.23

The SO₂ stream produced during the smelting step is sent to an acidplant for sulfuric acid production. The SO₂ content of the off gas inthis Example is 46%. Significant energy can be captured during sulfuricacid production, and this energy can be used to improve the overallenergy balance of the Looping Sulfide Oxidation process.

Slag Treatment

The slag produced in the electric furnace (3.0% Al₂O₃, 31.3% SiO₂, 15.5%CaO, 23.3% FeO, 13.6% Fe₂O₃, 13.2% Cu₂O) is transferred to an electricalfurnace at 1300° C. for slag treatment (Table 2). In this Example the3596.9 kg of slag is treated with 52 kg of carbon to reduce Fe₂O₃ andCu₂O. By reducing the trivalent iron, the solubility of copper in theslag is dramatically reduced. As a result, a copper melt is formed with97.7% of the copper recovered (417.1 kg melt, 98.997% Cu, 1.0% Fe)(FIGS. 10 and 11). The remaining slag contains only 0.35% Cu₂O and isfit for disposal as waste (melting temperature, 1070° C.; viscosity 1.7poise at 1300° C.) (FIGS. 12 and 13). The copper melt produced duringslag treatment is blended with the copper melt from the electric furnaceto produce a copper stream (7025.1 kg, 98.798% Cu, 0.062% Fe, 0.828% S,0.313% O) for treatment in the anode furnace.

The heat required to perform the slag treatment will be provided byelectrical heating via the electric furnace. Natural gas for combustionheating can also be provided via tuyères.

Downer Reoxidation Furnace

In a downer furnace, molten copper is atomized and oxidized in situ tofine. particulate CuO. Atomizing the molten copper minimizes masstransfer limitations between the molten copper and the oxygen and leadsto near 100% conversion to CuO. This highly exothermic reaction providessignificant potential for energy capture. It is understood that moltenCuO is highly corrosive, so following oxidation cool air is introducedto solidify the CuO. The CuO is thus cooled down to 800° C. before itexits as a fine particulate and is recycled back at temperature to thefirst furnace. Looping of this material in this system at temperatureand at high processing speed enhances the overall energy balance of theprocess.

The flue gases are sent to an air/air heat exchanger, where the reactionair for the downer furnace and anode furnace are preheated to 400° C. inorder to maximize the thermal efficiency. The flue gas is then sent to aboiler where a significant portion of the energy is captured as highpressure steam.

TABLE 3 Reoxidation Heat & Material Balance Temperature Amount, Amount,Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUT ReoxidationCopper 1300.000 93.105 5915.483 0.862 1284.49 1285.11 Cu 1300.000 93.0295911.598 0.660 1283.51 1283.51 Fe 1300.000 0.066 3.676 0.000 0.89 0.89O(g) 1300.000 0.009 0.144 0.202 0.07 0.69 S 1300.000 0.002 0.065 0.0000.02 0.02 Reaction Air 400.000 221.497 6390.255 4964.539 690.25 690.25O2(g) 400.000 46.514 1488.402 1042.553 150.10 150.10 N2(g) 400.000174.982 4901.853 3921.986 540.15 540.15 OUTPUT Copper Oxides 1243.85046.514 6661.065 1.109 1421.58 −502.52 Cu2O 1243.850 21.882 3131.1000.522 585.95 −438.94 Cu2O(1) 1243.850 24.600 3520.000 0.587 832.15−57.68 Cu2O*Fe2O3 1243.850 0.033 9.965 0.000 3.48 −5.90 Flue Gas1243.850 198.197 5644.748 4442.302 2161.41 2161.24 N2(g) 1243.850174.982 4901.853 3921.986 1895.51 1895.51 O2(g) 1243.850 23.212 742.764520.270 265.86 265.86 SO2(g) 1243.850 0.002 0.131 0.046 0.04 −0.13

TABLE 4 Quench Cooling of Reoxidation Products Temperature Amount,Amount, Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUTReoxidation- Quench Cooling Copper Oxides 1243.850 46.514 6661.065 1.1091421.58 −502.52 Cu2O 1243.850 21.882 3131.100 0.522 585.95 −438.94Cu2O(1) 1243.850 24.600 3520.000 0.587 832.15 −57.68 Cu2O*Fe2O3 1243.8500.033 9.965 0.000 3.48 −5.90 Flue Gas 1243.850 198.197 5644.748 4442.3022161.41 2161.24 N2(g) 1243.850 174.982 4901.853 3921.986 1895.51 1895.51O2(g) 1243.850 23.212 742.764 520.270 265.86 265.86 SO2(g) 1243.8500.002 0.131 0.046 0.04 −0.13 Cooling Air 25.000 528.256 15240.36611840.122 0.00 0.00 N2(g) 25.000 417.322 11690.618 9353.697 0.00 0.00O2(g) 25.000 110.934 3549.748 2486.426 0.00 0.00 OUTPUT Copper Oxides800.000 93.029 7405.274 1.172 1026.09 −3046.62 CuO 800.000 92.9967397.400 1.172 1024.68 −3039.20 CuO*Fe2O3 800.000 0.033 7.874 0.000 1.41−7.42 Flue Gas 800.000 703.196 20140.911 15761.146 4705.52 4705.35 N2(g)800.000 592.305 16592.471 13275.683 3927.10 3927.10 O2(g) 800.000110.889 3548.310 2485.418 778.39 778.39 SO2(g) 800.000 0.002 0.131 0.0460.02 −0.15

In this Example, 7400 kg of CuO are required in the electric furnace. Assuch, 5911.1 kg of molten Cu must be oxidized in the downer reoxidationfurnace; the remaining 1020.5 kg of Cu can be sent to electrolysis forfinal purification (Tables 3 and 4). In the downer reoxidation furnace asignificant excess of air will be used to ensure complete reoxidation.

Energy is captured during this step by using the flue gases from thereoxidation furnace to (1) preheat the oxidation air and (2) producehigh pressure steam in a boiler after preheating.

Energy Balance

The two primary energy producing steps in the Looping Sulfide Oxidationprocess are the sulfuric acid production in the acid plant and thereoxidation of the Cu to CuO before it is looped back to the electricfurnace. The acid plant per se, is outside the scope of this invention;however, as it is known to those skilled in the art, state-of-the-artprocesses like the Lurec® process have been shown to capture significantportions of the total energy available during sulfuric acidproduction¹⁴. On this basis, we have evaluated the energy balance of theLooping Sulfide Oxidation process relative to conventional copperprocessing. ¹⁴ Daum K H, The Lurec® Process—Key to Economic Smelter AcidPlant Operation, in The Southern African Institute of Mining andMetallurgy Sulfur and Sulfuric Acid Conference 2009, 1-22.

During conventional copper processing, the only major energy producingstep is the acid production. It is estimated that the theoretical totalamount of energy that can be produced during this step is 54.7 Wh permole of CuFeS₂ processed.

2SO₂+O₂→2SO₃ ΔH_(600° C.)=−54.7 Wh  (6)

In this analysis, production of sulfuric acid is estimated to result inthe production of 2462 kg of high pressure steam (100 bar, 350° C.) per1000 kg of Cu produced during heat capture in boilers and coolingjackets (Tables 5 and 6).

TABLE 5 Acid Plant Boiler Heat & Material Balance Temperature Amount,Amount, Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUT AcidPlant System- Boiler 1 Gas from Smelting 1300.000 73.419 3407.2211645.574 1143.46 −2551.06 SO2(g) 1300.000 33.772 2163.415 756.960 634.42−2150.05 CO2(g) 1300.000 8.326 366.412 186.609 152.67 −757.38 N2(g)1300.000 31.320 877.394 702.005 356.37 356.37 Gas from Anode 1200.00050.173 1437.869 1124.550 580.51 −561.04 Furnace N2(g) 1200.000 35.269987.991 790.495 367.09 367.09 SO2(g) 1200.000 1.810 115.964 40.575 31.08−118.18 O2(g) 1200.000 0.003 0.082 0.057 0.03 0.03 NO(g) 1200.000 0.0010.026 0.019 0.01 0.03 SO3(g) 1200.000 0.001 0.052 0.015 0.02 −0.06H2O(g) 1200.000 8.273 149.040 185.428 108.08 −447.65 CO2(g) 1200.0003.849 169.393 86.270 64.33 −356.39 CO(g) 1200.000 0.514 14.407 11.5285.41 −10.38 H2(g) 1200.000 0.453 0.914 10.163 4.46 4.46 Cooling Water25.000 89.461 1611.655 1.617 0.00 −7098.80 H2O(100 barl) 25.000 89.4611611.655 1.617 0.00 −7098.80 OUTPUT Gas to Scrubbing 400.000 123.5914845.090 2770.123 465.76 −4370.31 SO2(g) 400.000 35.583 2279.379 797.534171.29 −2762.43 CO2(g) 400.000 12.175 535.805 272.879 55.58 1275.20N2(g) 400.000 66.589 1865.385 1492.500 205.55 205.55 O2(g) 400.000 0.0030.082 0.057 0.01 0.01 NO(g) 400.000 0.001 0.026 0.019 0.00 0.02 SO3(g)400.000 0.001 0.052 0.015 0.00 −0.07 H2O(g) 400.000 8.273 149.040185.428 30.33 −525.40 CO(g) 400.000 0.514 14.407 11.528 1.62 −14.17H2(g) 400.000 0.453 0.914 10.163 1.37 1.37 High Pressure 350.000 89.4611611.655 2005.140 485.79 −5840.59 Steam H2O(100 barg) 350.000 89.4611611.655 2005.140 485.79 −5840.59

TABLE 6 Catalyst Bed Heat & Material Balance Temperature Amount, Amount,Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUT Catalyst BedPost-Scrub Gas Stream 80.000 114.346 4680.570 2562.913 59.08 −4205.41SO2(g) 80.000 35.583 2279.379 797.534 22.29 −2911.43 N2(g) 80.000 66.5891865.385 1492.500 29.65 29.65 CO2(g) 80.000 12.175 535.805 272.879 7.14−1323.64 Reaction it 25.000 93.193 2688.636 2088.781 0.00 0.00 O2(g)25.000 19.570 626.230 438.644 0.00 0.00 N2(g) 25.000 73.622 2062.4061650.137 0.00 0.00 Cooling Water 25.000 47.217 850.633 0.853 0.00−3746.75 H2O(100 barl) 25.000 47.217 850.633 0.853 0.00 −3746.75 OUTPUTCatalyzed Gas 225.000 189.748 7369.206 4252.926 373.05 −4869.49 SO3(g)225.000 35.583 2848.680 797.534 114.14 −3797.62 N2(g) 225.000 140.2113927.791 3142.636 228.08 228.08 CO2(g) 225.000 12.175 535.805 272.87927.84 −1302.94 O2(g) 225.000 1.779 56.930 39.877 2.98 2.98 High PressureSteam 350.000 47.217 850.633 1058.314 256.40 −3082.67 H2O(100 barg)350.000 47.217 850.633 1058.314 256.40 −3082.67

Therefore, with all other factors being equal, conventional copperprocessing and Looping Sulfide Oxidation processing would theoreticallyproduce equal amounts of energy during sulfuric acid production.However, as the Lurec® process states, the higher the strength of theSO₂ stream, the greater the energy production; therefore, it can beexpected that, in practice, the Looping Sulfide Oxidation process wouldactually produce more energy than the conventional process due to itshigh strength SO₂ stream. However, if equal energy production is assumedin the acid plant, the only major differentiating factor in energyproduction will be during the reoxidation of the copper to CuO, whichthe conventional process does not perform. During reoxidation, theamount of high pressure steam (100 bar, 350° C.) that is estimated to beproduced is 4049 kg per 1000 kg of Cu produced (Tables 7 and 8).

TABLE 7 Reoxidation Reaction Air Preheater Heat & Material BalanceTemperature Amount, Amount, Amount, Latent Total © kmol kg Nm3 H, kWh H,kWh INPUT Reoxidation Heat Recovery-Air Preheater New Reaction Air25.000 266.141 7678.264 5965.184 0.00 0.00 O2(g) 25.000 55.890 1788.4021252.689 0.00 0.00 N2(g) 25.000 210.252 5889.862 4712.495 0.00 0.00Reoxidation Flue 800.000 703.196 20140.911 15761.146 4705.52 4705.35Gases N2(g) 800.000 592.305 16592.471 13275.683 3927.10 3927.10 O2(g)800.000 110.889 3548.310 2485.418 778.39 778.39 SO2(g) 800.000 0.0020.131 0.046 0.02 −0.15 OUTPUT New Reaction Air 400.000 266.141 7678.2645965.184 829.38 829.38 O2(g) 400.000 55.890 1788.402 1252.689 180.35180.35 N2(g) 400.000 210.252 5889.862 4712.495 649.02 649.02 ReoxidationFlue 671.626 703.194 20140.780 15761.101 3875.97 3875.97 Gases N2(g)671.626 592.305 16592.471 13275.683 3235.65 3235.65 O2(g) 671.626110.889 3548.310 2485.418 640.32 640.32 SO2(g) 671.626 0.000 0.000 0.0000.00 0.00

TABLE 8 Reoxidation Boiler Heat & Material Balance Temperatur Amount,Amount, Amount, Latent Total © kmol kg Nm3 H, kWh H, kWh INPUTReoxidation Heat Recovery-Boiler Reoxidation Flue 671.626 703.19420140.780 15761.101 3875.97 3875.97 Gases N2(g) 671.626 592.30516592.471 13275.683 3235.65 3235.65 O2(g) 671.626 110.889 3548.3102485.418 640.32 640.32 SO2(g) 671.626 0.000 0.000 0.000 0.00 0.00Cooling Water 25.000 224.748 4048.881 4.061 0.00 −17833.96 H2O(100 barl)25.000 224.748 4048.881 4.061 0.00 −17833.96 OUTPUT Reoxidation Flue150.000 703.194 20140.780 15761.101 715.05 715.05 Gases N2(g) 150.000592.305 16592.471 13275.683 600.31 600.31 O2(g) 150.000 110.889 3548.3102485.418 114.73 114.73 SO2(g) 150.000 0.000 0.000 0.000 0.00 0.00 HighPressure 350.000 224.748 4048.881 5037.413 1220.43 −14673.04 SteamH2O(100 barg) 350.000 224.748 4048.881 5037.413 1220.43 −14673.04

Taking into consideration the total estimated energy output duringLooping Sulfide Oxidation, the amount of energy available for captureduring the reoxidation of the molten copper is approximately 1.64 timesgreater than the amount available for capture during sulfuric acidproduction alone. This comparison is vital because during conventionalprocessing, significant energy consumptions and productions have beenobserved at different processing facilities¹⁵. Therefore, on the basisof potential energy available for capture, the Looping Sulfide Oxidationprocess provides significant improvements over the conventionaltechnology; the increased energy production drastically mitigates thenet energy consumption during copper processing. ¹⁵Coursol P, Mackey PJ, and Diaz C M (2010) Energy Consumption in Copper Sulphide Smelting,in Proceedings of Copper 2010, 1-22.

Example 2

Using the same feed conditions and smelting furnace parameters as thosepresented in Example 1, the slag produced in the smelting furnace can betreated in the slag treatment furnace by sulfidation. Duringsulfidation, iron pyrite (FeS₂) is added to the molten slag to sulfidizethe copper, causing it to separate out of the slag into a copper matte(FIGS. 14 and 15). In this scheme, the copper recovery from the slagranges from 99 to 96% in the temperature range of 1200-1400° C. The slaghas a melting temperature of 1120° C. and a viscosity of 0.709 poise at1300° C. The treated slag is fit for disposal as waste. The coppermatte, which is now rich in copper sulfide, must be processed in thesmelting furnace again before the copper can be sent to the anodefurnace as blister copper.

Example 3

Copper sulfide concentrate (CuFeS₂) is smelted with CuCO₃ to producecopper metal, iron oxide slag, and rich SO₂ off gas (FIGS. 16-18). Insuch a reaction, 3000 kg of CuFeS₂ (with 173.4 kg of FeS₂ and 294.8 kgof CaAl₂Si₂O₈) is reacted with 11500 kg of CuCO₃ and 1000 kg of SiO₂ and500 kg of CaO between 1200° C. and 1400° C. The products of thisreaction will include an off gas that is comprised mainly of CO₂ and SO₂(FIG. 19). At 1300° C., 6654 kg of molten Cu will be produced containing0.30% S, 0.21% O and 0.0028% Fe. The 3490 kg of slag produced contains10.7% Cu₂O.

Example 4

Copper sulfide concentrate (CuFeS₂) is smelted with CuSO₄ to producecopper metal, iron oxide slag and rich SO₂ off gas (FIG. 20-22). In sucha reaction, 3000 kg CuFeS₂ (with 173.4 kg FeS₂, 294.8 kg CaAl₂Si₂O₈) isreacted with 7423 kg CuSO₄ and 1000 kg SiO₂ and 500 kg CaO between 1200°C. and 1400° C. The products of this reaction will include an off gasthat is comprised of SO₂ that is diluted with any combustion gases orinert gases. At 1300° C., 3658 kg of molten copper will be producedcontaining 1.1% S, 0.33% O and 0.0025% Fe. The 3559 kg of slag producedcontains 12.3% Cu₂O.

It will now be apparent to those skilled in the art that otherembodiments, improvements, details, and uses can be made consistent withthe letter and spirit of the foregoing disclosure and within the scopeof this patent, which is limited only by the following claims, construedin accordance with the patent law, including the doctrine ofequivalents.

What is claimed is:
 1. A method for production of copper comprising: (a)providing, (1) a copper sulfide concentrate product of mineral refiningcomprising copper and iron metal values including sulfides thereof and(2) one or more copper oxides, to a molten slag wherein they react witheach other for smelting desulfurization, (b) the copper oxides beingprovided to the slag in stoichiometric or in slight excess (up to about20%) of stoichiometric ratios (c) agitating the molten slag via theinjection of the gaseous products of fossil fuel combustion and/or achemically inert gas (d) thereby oxidizing the sulfide concentrate tomolten blister copper by conversion essentially with one or more copperoxides in a one step, molten bath operation to produce: (1) moltenblister copper, (2) iron oxide slag, and (3) highly concentrated SO₂ offgas, (e) treating the blister copper to reduce the iron content andoxidize residual sulfur therein, and to prepare it for eitherelectrolysis or reoxidation.
 2. The method of claim 1, wherein the ironoxide slag is treated in a slag treatment furnace by carbon (as coaland/or natural gas) reduction or sulfur oxidation to recover furthercopper metal and, wherein the recovered copper metal is provided to ananode furnace and treated therein.
 3. The method of claim 1, wherein theiron oxide slag is treated via sulfidation with iron pyrite to produce amatte and providing the resulting copper matte to a smelting furnace forconversion into molten blister copper.
 4. The method of claim 1, whereinthe SO₂ rich off gas is provided to one or more of plants selected fromthe group consisting of a handling plant for sulfuric acid production, agypsum production plant, and a sulfur dioxide liquefaction plant.
 5. Themethod of claim 1 wherein the copper sulfide concentrate oxidation isperformed with Cu₂O.
 6. The method of claim 5, wherein the stoichiometryof the reaction in the smelting furnace, in which the copper feed isconverted to metallic copper, a slag, and an SO₂ off gas, is defined asthe amount of Cu₂O required to completely (1) convert the coppercontained in the feed to metallic copper, (2) oxidize any iron in thefeed to FeO and/or Fe₂O₃, which report to the slag, and (3) oxidize anysulfur in the feed to SO₂ and substantially maintained in providing thesulfide and oxide.
 7. The method of claim 5, wherein a flux material isprovided and the copper concentrate, flux and Cu₂O materials are fedinto the molten slag where they react before separating to theirrespective phases.
 8. The method of claim 1, wherein the copper sulfideconcentrate oxidation is performed with CuO.
 9. The method of claim 8,wherein the stoichiometry of the reaction in the smelting furnace, inwhich the copper feed is converted to metallic copper, a slag, and anSO₂ off gas, is defined as the amount of copper oxides CuO required tocompletely (1) convert the copper contained in the feed to metalliccopper, (2) oxidize any iron in the feed to FeO and/or Fe₂O₃, whichreport to the slag, and (3) oxidize any sulfur in the feed to SO₂ andsubstantially maintained in providing the sulfide and oxide.
 10. Themethod of claim 8, wherein a flux material is provided and the copperconcentrate, flux and CuO are fed into the molten slag where they reactbefore separating to molten slag and blister copper.
 11. The method ofclaim 1, wherein the temperature of the smelting furnace, wherein theoxidation of the copper concentrate is performed, is 1100-1400° C. 12.The method of claim 1 wherein a chemically inert gas is injected intothe molten slag formed during the oxidation of the copper concentrate topromote chemical reaction.
 13. The method of claim 12, wherein thechemically inert gas is N₂.
 14. The method of claim 12, wherein thechemically inert gas comprises one or more combustion products of fossilfuel.
 15. The method of claim 12, wherein the chemically inert gas is anoble gas.
 16. The method of claim 12, wherein the furnace for theoxidation of the copper concentrate is an electric furnace with tuyèresto blow the chemically inert gas into the molten slag.
 17. The method ofclaim 1, wherein the furnace for the oxidation of the copper concentrateis an induction furnace.
 18. The method of claim 17, wherein theinduction heating is operated to facilitate mixing in the slag.
 19. Themethod of claim 1, wherein the sulfur in the copper sulfide concentrateis oxidized to produce an oxidized sulfur gas with a content of 20-100%SO₂ without the use of oxygen-enriched air.
 20. The method of claim 1,wherein the sulfur content in the molten copper is reduced to below 1%,the iron content in the molten copper is reduced to below 0.3% and theoxygen content in the molten copper is reduced to below 0.6%.
 21. Themethod of claim 20, wherein the sulfur content in the molten copper isreduced to below 0.9%, and the iron content in the molten copper isreduced to below 0.002%.
 22. The method of claim 1, wherein the slag istreated to recover copper.
 23. The method of claim 22, wherein thefurnace used to treat the slag is an electric furnace.
 24. The method ofclaim 22, wherein the slag treatment is performed with carbon and theresulting copper melt is fed to an anode furnace.
 25. The method ofclaim 22, wherein the slag treatment is performed with sulfur or sulfurcontaining compounds and the resulting copper matte is fed to thesmelting furnace.
 26. The method of claim 22, wherein the residualcopper content in the treated slag is below 0.5% and the total copperrecovery from the slag exceeds 92%.
 27. The method of claim 1, whereincopper is reoxidized with air to produce the required amount of copperoxide for use in the smelting-desulfurization step.
 28. The method ofclaim 27, wherein at least 80% CuO relative to capacity of the copper tobe reoxidized is produced.
 29. The method of claim 1, wherein moltencopper is atomized to molten droplets and reoxidized in a vertical,flash or downer furnace.
 30. The method of claim 1, wherein the moltencopper is atomized to solid copper powder before it is reoxidized. 31.The method of claim 30, wherein the atomized solid copper powder isreoxidized in a kiln, rotary kiln, fluid bed, flash, downer, shaft, ormultiple hearth furnace.
 32. The method of claim 1 wherein scrap coppermetal and/or alloys are oxidized in a reoxidation furnace to producecopper oxide(s), which can be directed to the smelting furnace.
 33. Themethod of claim 1 wherein copper oxides are provided from one or moreexternal sources selected from the group consisting of pigments, spentcatalysts, battery components, and one or more of the minerals such asmalachite, azurite, cuprite, chrysocolla, blue vitriol, antlerite,brochantite.
 34. The method of claim 1 wherein the external sources ofcopper oxygen carriers are provided that are selected from the groupconsisting of sulfates, carbonates, hydroxides, and one or more mineralssuch as malachite, azurite, cuprite, chrysocolla, blue vitriol,antlerite and brochantite.
 35. The method of claim 4 wherein energy isproduced and captured by the further step of producing sulfuric acidfrom the rich SO₂ off gas from the smelting furnace.
 36. The method ofclaim 35, wherein reoxidation of copper to copper oxides produces morethan 1.5 times energy than is produced in a sulfuric acid plant handlingthe SO₂ off gas from the smelting furnace.
 37. A method for productionof copper comprising the following steps: (a) copper sulfide concentrateand one or more copper oxides are fed into a molten slag wherein theyreact with each other (b) the molten slag is agitated via the injectionof the gaseous products of fossil fuel combustion or a chemically inertgas with air, whose oxygen further reacts with the copper sulfideconcentrate and one or more copper oxides (c) the total of copper oxidesand oxygen from air are fed in stoichiometric or in slight excess (up toabout 20%) of stoichiometric ratios (d) oxidation of copper sulfideconcentrate to molten blister copper is carried out by conversion with amixture of copper oxides and oxygen in air in a one step, molten bathoperation to produce (1) molten blister copper, (2) iron oxide slag, and(3) highly concentrated SO₂ off gas, (e) the blister copper is treatedto reduce the iron content and oxidize residual sulfur, and prepare itfor either electrolysis or reoxidation.
 38. The method of claim 37,wherein the iron oxide slag is treated in a slag treatment furnace bycarbon (as coal and/or natural gas) reduction and wherein the moltenblister copper is provided to an anode furnace.
 39. The method of claim37, wherein the iron oxide slag is treated via sulfidation with ironpyrite and wherein the resulting copper matte is provided to a smeltingfurnace for oxidation.
 40. The method of claim 37, wherein the SO₂ richoff gas is sent to a handling plant for sulfuric acid production, gypsumproduction and/or sulfur dioxide liquefaction.
 41. The method of claim37, wherein the process by which the copper sulfide concentrate isoxidized is performed with CuO and O₂ in air without oxygen-enrichment.42. The method of claim 37, wherein the process by which the coppersulfide concentrate is oxidized is performed with Cu₂O and O₂ in airwithout oxygen-enrichment.
 43. The method of claim 37, wherein theoxygen used during smelting, in processing of the SO₂ rich off gas, andreoxidation is not enriched to increase the O₂ content relative to N₂ inair (with about 21% O₂, 79% N₂ proportions).
 44. The method of claim 37,wherein the stoichiometry of the reaction in the smelting furnace, inwhich the copper feed is converted to blister copper, a slag, and an SO₂off gas, is defined as the amount of copper oxides (CuO and/or Cu₂O) andO₂ (from natural air) required to completely (1) convert the coppercontained in the feed to metallic copper, (2) oxidize any iron in thefeed to FeO and/or Fe₂O₃, which report to the slag, and (3) oxidize anysulfur in the feed to SO₂ and is substantially maintained in providingsulfide and oxide.
 45. The method of claim 44, wherein the relativeratios of CuO and/or Cu₂O to O₂ (from natural air) may vary to any limitprovided the total minimum stoichiometry is met.
 46. The method of claim37 wherein the oxidizing agent for the oxidizing step comprises CuO andwherein the method for feeding the copper concentrate, flux and CuO issuch that both materials are fed into the molten slag where they reactbefore separating to molten slag and blister copper.
 47. The method ofclaim 37, wherein the oxidizing agent for the oxidizing step comprisesCu₂O and wherein the method for feeding the copper concentrate, flux andCu₂O is such that both materials are fed into the molten slag where theyreact before separating to molten slag and blister copper.
 48. Themethod of claim 37, wherein the temperature of the smelting furnace,wherein the oxidation of the copper concentrate is performed is1100-1400° C.
 49. The method of claim 37, wherein the air is injectedinto the molten slag formed during the oxidation of the copperconcentrate to promote chemical reaction.
 50. The method of claim 37,wherein a chemically inert gas is injected into the molten slag formedduring the oxidation of the copper concentrate to promote chemicalreaction.
 51. The method of claim 50, wherein the chemically inert gasis N₂.
 52. The method of claim 50, wherein the chemically inert gas isthe combustion products of a fossil fuel.
 53. The method of claim 50,wherein the chemically inert gas is a noble gas.
 54. The method of claim50, wherein the furnace for the oxidation of the copper concentrate isan electric furnace with tuyères to blow the chemically inert gas intothe molten slag.
 55. The method of claim 50, wherein the furnace for theoxidation of the copper concentrate is an induction furnace.
 56. Themethod of claim 55, wherein the induction heating facilitates mixing inthe slag.
 57. The method of claim 37, wherein the sulfur in the coppersulfide concentrate is oxidized to produce an oxidized sulfur gas with acontent of 20-100% SO₂ without the use of oxygen-enriched air.
 58. Themethod of claim 37, wherein the sulfur content in the molten copper isreduced to below 1%, the iron content in the molten copper is reduced tobelow 0.3%. The oxygen content in the molten copper is below 0.6%. 59.The method of claim 37, wherein the sulfur content in the molten copperis reduced to below 0.9% and the iron content in the molten copper isreduced to below 0.002%.
 60. The method of claim 38, wherein the slag istreated to recover copper.
 61. The method of claim 60, wherein thefurnace used to treat the slag is an electric furnace.
 62. The method ofclaim 60, wherein the slag treatment is performed with carbon and theresulting blister copper is provided to the anode furnace.
 63. Themethod of claim 60, wherein the slag treatment is performed with sulfuror sulfur containing compounds and the resulting copper matte isprovided to the smelting furnace.
 64. The method of claim 60, whereinthe residual copper content in the treated slag is reduced below 0.5%and the total copper recovery from the slag exceeds 92%.
 65. The methodof claim 37, wherein copper is reoxidized with air to produce therequired amount of copper oxide for use in the smelting-desulfurizationstep.
 66. The method of claim 65, wherein at least 80% CuO relative tocapacity of the copper to be reoxidized is produced.
 67. The method ofclaim 65, wherein molten copper is atomized to molten droplets andreoxidized in a vertical, flash or downer furnace.
 68. The method ofclaim 65, wherein the molten copper is atomized to solid copper powderbefore it is reoxidized.
 69. The method of claim 68, wherein theatomized solid copper powder is reoxidized in a kiln, rotary kiln, fluidbed, flash, downer, shaft, or multiple hearth furnace.